CA2076025A1 - Recovery of metal values from zinc plant residues - Google Patents
Recovery of metal values from zinc plant residuesInfo
- Publication number
- CA2076025A1 CA2076025A1 CA 2076025 CA2076025A CA2076025A1 CA 2076025 A1 CA2076025 A1 CA 2076025A1 CA 2076025 CA2076025 CA 2076025 CA 2076025 A CA2076025 A CA 2076025A CA 2076025 A1 CA2076025 A1 CA 2076025A1
- Authority
- CA
- Canada
- Prior art keywords
- zinc
- copper
- iron
- flotation
- leach
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Abandoned
Links
Classifications
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B13/00—Obtaining lead
- C22B13/04—Obtaining lead by wet processes
- C22B13/045—Recovery from waste materials
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B11/00—Obtaining noble metals
- C22B11/04—Obtaining noble metals by wet processes
- C22B11/042—Recovery of noble metals from waste materials
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B15/00—Obtaining copper
- C22B15/0063—Hydrometallurgy
- C22B15/0065—Leaching or slurrying
- C22B15/0067—Leaching or slurrying with acids or salts thereof
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/20—Obtaining zinc otherwise than by distilling
- C22B19/22—Obtaining zinc otherwise than by distilling with leaching with acids
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/30—Obtaining zinc or zinc oxide from metallic residues or scraps
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/006—Wet processes
- C22B7/007—Wet processes by acid leaching
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
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- Engineering & Computer Science (AREA)
- Chemical & Material Sciences (AREA)
- Manufacturing & Machinery (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Metallurgy (AREA)
- Organic Chemistry (AREA)
- Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Geology (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
A process is disclosed for recovering zinc, lead, copper and precious metals from zinc plant residue, said process comprising leaching the residue with return zinc spent electrolyte, neutralizing residual acid and reducing ferric iron in the solution by addition of zinc sulphide concentrate in the presence of a limited quantity of oxygen, flotation of the resulting slurry to separate unreacted zinc sulphide, treatment of the flotation tailings with sulphur dioxide and elemental sulphur to further leach iron, zinc and impurity elements and precipitate copper, flotation of the resulting slurry to separate a copper sulphide concentrate, thickening, filtering and washing of the flotation tailings followed by addition of lime and sodium sulphide to activate lead sulphate and flotation of a lead concentrate from the residue. Iron and impurity elements are precipitated from the copper flotation tailings thickener overflow solution by addition of zinc hydroxide sludge, lime and oxygen to produce a high strength, iron free zinc sulphate solution.
Description
2~7~2~
- I - PCT/C.~90/00442 RECOVERY OF METAL VALUES FROM Z INC PLANT RESIDUES
~IELD 0~ T~E I~ENTION
This lnventlon ~elates to the recovery of metal values from z~nc plant r~sidues and, more particularly, relates to the separation of zinc, lead. copper and precious metal values from zinc plant residues in a form amenable to the recovery of these metal values.
BACXGROUND OF THE INVENTION
Residues produced in the treatment of zinc concentrates by conventional roast-leach-electrowinning processes may contain significant ~uantlties of zinc, but the recovery of this zine may be hampered by the presence of interfering elemen~s such a~ iron or undesirable impurities such as arsenic, aermanium or fluoride. It has been well known for many years that trea~ment of zlnc ferrite residues in a hot acid leach will solubilize the zinc. Howe~er. treatment of these residues to recover the contained zinc was not economic until the 1960's when the jarosite and goethite proeesses w-re developed for handllng the iron dissolved fro~ the ferrite residues. In certain plants, the inplementatlon of these processes has allowed ~or treatment of stockpiled residues which are fed to the hot acid leach at a slow rate along with current material.
Not all zinc process residues are amenable to such a treatment, however, owing to the nature of the zinc species in the residue, which may be refractory to the hot acid leach process, or to the presence of other elements, wh~ch either interfere with the recovery of zinc or must be recovered in addition to the zinc. For exacple, the zinc plant residue may contain significant quantities of metal values other than zinc, such as lead, copper or precious metals, but the low grade of the residue in respect of these elements may preclude their economic recovery on an individual basis.
U.S. Patent No. 4,572,822 issued February 25, 1986 to Abe and Tanaka discloses a process for recovering copper from industrial residues involving a reducing leach of the residue in sulphur dioxide atmosphere to dissolve oxidtc - '' ' : .
P~T/CA90~0442 2~76025 - 2 cop~er compounds and the addition of elemental sulphur to preclpitate copper from the solutlon as cop~er sulphide.
Althouah this process allows for the recovery of copper found in the industrial residue in both sulphidic and oxidic forms as a single copper sulphide product, the process does not allow for the presence of zinc as zinc sulphide in the industrial residue. Zinc sulphlde, if present, would report along with the copper sulphide in a subsequent step to separate the copper sulphide from the industrial residue and would render the copper sulphide product of little or no value if the zinc sulphide were present in an amount equal to or greater than the amount of the copper sulphide. '!
In accordance with the process of the present lnvention, copper sulphide product from an industrial residue is recoverable substantially free of zinc.
U.S. ~atent No. 4,676,828 ~ssued June 30, 1987 to Andre discloses a process for the extraction of zinc from zinc sulphide concentrate including leaching with a dilute aqueous solution of sulphuric acid under atmospheric pressure. This process, however, relates to the extractlon of zinc and copper from sulphurous zinciferous materials, and does not permit the separation and recovery of copper and zinc from industrial residues which contain a large proportion of oxidic compounds.
The flotation of lead sulphate and the precipitation of iron from a zinc sulphate solution are discussed in Papers by Fuerstenau et al, ''The Surface Characteristics and Flotation Behavior of Anglesite and Cerussite", Int. J. of Min. Proc., 1987, 20, 73-85: Andre and Masson, "The Goethite Process in Retreating Zinc Leaching Residues", presented at the 102nd AIME Annual Meeting, 1973, Chicago; 30xall and James, "Experience with the Goethite Process at National Zinc"; and in J.E.
Dutrizac and A.J. Monhemius ~Ed.), Iron Control in Hydrometallurgy, 1986, Ellis Horwood, Chichester, 676-686.
.
: .
~O 91/091 ~h PCT/CA90/0 ~ 2 ~ 3 ~ 20~25 None of these Papers discloses or suagesls the comblnatlon of steps and parameters of the process c' the present lnvention for the recovery of metal values from industrial residues The process of the invention permits separation of zinc. copper, lead and precious metals from zinc plant resldues in a readily recoverable form and surprisingly allows for recovery of metal values present in the head material in two or more substantially different mineral forms. Zinc or copper, for example. may be recovered in high yields whether they are present as sulphidic or oxidic compounds, or both.
SUMMARY OF T~E INVENTION
In accordance with the process of the present inventior.. -inc plant residue is treated in both hot acid leaching and reducing leaching stages, with a neutralization and reduction stage and a flotation stage between the two leaching stages. Copper sulphide in the residue is first reacted with ferric iron in solution in the hot acid leach stage to oxidize the copper sulphide and to brina the copper conte~t of this sulphide into the solution, Zinc sulphide in the residue is then recovered by flotation separation prio- to leaching the remaining oxidic copper compounds and precipitating copper from the solution in the reducing leach stage with sulphur dioxide and elemental sulphur. The copper sulphide product that is recovered is substantially free of zinc. As a result of this combination of steps, both copper and zinc may be present in oxidic and sulphidic forms in the feed to the process and be recovered separately to hi~h grade concentrates or to a high stren~th solution.
More particularly, the process of the invention relates to the recovery of zinc, lead, copper and precious metals from zinc plant residue containing ferrites compris-ina the steps of leaching said zinc plant residues with return spent electrolyte containina ~ S04 in an amount effective to dissolve the ferrites and to maintain at least 50 g/L ~ S04 in a hot acid leach z: a temperature in the range of . . ~ ,. ' , .
.~ ~ .. . .
., .
~ ' .
,.:
~ 0 91/09146 PCT/CA90/1)0~142 2 ~ 7 6 ~ 2 ~
70O to 100C and at atmospherlc pressure to partlally dissolve zinc, copper. iron and impurlty elements and to e~sentially leach sulphide copper; treating the resulting 10ach slurry with zinc concentrate under oxidizing ccnditions at atmospheric pressure at a temperature in the range of 70 to 100C to consume excess acid from said hot acid leach and to increase the concentration of zinc in the leach solution and continulng said treatment with zinc concentrate under reducing conditions to reduce ferric iron in solution to ferrous iron; recovering excess zinc concentrate and elemental sulphur produced in the reaction of zinc concentrate with acid and ferric iron by zinc flotation as a flotation concentrate: subjecting ~inc flotation tailings to a reducing leach in the presence of gaseous sulphur dioxide and elemental sulphur at a temperature in the range of 700 to 120C with a sulphur dioxide overpressure of at least 30 kPa to extract zinc, copper, iron and impurity elements, to reprecipitate copper as coppor sulphlde, and to convert load in jarosite to lead sulphate: recovering copper sulphlde by copper flotation as a ~lotation concentrate: subjecting the copper flotation tailings to a liguid-solid separation;
reco~ering lead sulphate as a flotation concentrate by lead flotation of the separated solids from the liquid-solid separation of the copper flotation tailings; and recovering and treating the reducing leach solution from the liquid-solid separation for the recover~ of contained zinc values.
The process preferably includes adjusting the pH of the reducing leach solution from the liquid-solid separation to about 3.5 to 4.0 and oxidizing the largest part of contained ferrous iron to ferric iron in an oxidizing atmosphere at a temperature in the range of 700 to 100C. preferably about 85oC, at atmospheric pressure for precipitation of iron as ferric hydroxide or hydrated ferric oxide with impurity elements, and separating the " : ' ' '' - ~ ' ~ ' ., ' ; ' , ' -, ~091/09116 PCT/CA90/00~42 - 5 _ 2976~2~
resldual solutlon: and adjustin~ the pH of the recovered reSldUal sOlUtlOn tO about 5.0, oxidizing the contained iron to ferric iron in an oxidizing atmosDhere at a temperature in the range of 700 ro 100C, preferably about 8~oC. at atmospheric Dressure for precipitation of iron as ferric hydroxide or hydrated ferric oxide, and separating the residual solution for recovery of contained zinc values.
DESCRIPTION OF THE ~RAWING
The process of the invention will be described with reference to the flow sheet of the accompanying drawing.
DESCRIPTION OF THE PREFERRED EMBODIMENT
Zinc plant residue containing zinc, lead, copper and precious metal values is contacted with return zinc spent electrolyte at atmospheric pressure and a temperature of at least 70C, preferably in the range of 700 to 100C, and more preferably, 850 to 950C to partially dissolve zinc, copper, iron and impurity elements. It is of particular importance to the recovery of copper that sulphide copper in the zinc pla~t residue is leached ln this stcp, The resultins slurry i9 treated with zinc concentrate under oxygen sparging conditions at atmospheric pressure and a temperature of at least 700C or more preferably 850 to 950C, to consume the excess of acid from the previous step and to increase the concentration of zinc in the solution. Additional zinc concentrate is fed to the slurry in the absence of oxygen, but at otherwise similar temperature and pressure, to reduce ferric iron in solution to the ferrous state. Excess zinc concentrate and elemental sulphur produced in .the reaction of zinc concentrate with acid and ferric iron is recovered by flotation, this flotation concentrate being suitable as feed to a zinc pressure leach autoclave or a roaster.
The flotation tailings is treated with sulphur dioxide and elemental sulphur at a temperature of at least 700C. preferablY in the range of 70 to 120C, and more .. :, ' ' ' :. . , .. ,. - ~, . .
' ' '"'" ' . ' :... ' ~, . , :
~091/09~6 PCT/CA90/n~2 207~02~ - 6 -~rererablv ln t~e ranae sOo ~o 100C and wlth an overpressure of sulphur dioxide of at least 30 kPa.
pre~erably at about 100 kPa. to further extract zlnc.
copper. lron and impurity elements. to repreclpitate copper as copper sulphide. and to convert lead in jarosite to lead sulphate. The slurry resulting from this reducina leach is sub~ected to a second flotation step to separate a copper sulphide concentrate, this flotation concentrate being suitable as feed to a copper smelter. After liq~id-solid separation and adjustment to between pH 8.5 and p~ 10.0, more prefrably pH 9.S, the flotation tailings solids are sulphidized for recovery of a lead sulphate flotatlon concentrate. which is suitable as feed to a lead smelter.
The lead flotation tailings, comprised of aypsum and other gangue minerals, is stoc~piled.
The leach solution separated from the copper flotation tailings solids, containing the largest part of the zinc in the head material, is treated with lime or zinc hydroxide sludge in a two stage process for precipitation of iron and impurity elements from the s~lution. Both stages are maintained at a temperature between 70 and 100C, more preferably at 85oC. The solution is maintained at about pH 4 or less in the first stage to limit the coprecipitation of zinc. After liquid-solid separation, the iron cake is impounded and the solution sent to a second stage of iron removal at pH 5. The resulting slurry is clarified and the solids recycled to the first stage of the iron removal step for recovery of precipitated zinc.
Zinc may be recoverd from the low iron zinc sulphate solution using conventional purification and electrowinning techniques, or the solution may be integrated with the existing zinc plant. Precious metal values in the oxide residue head material are distributed among the three flotation concentrates in the process, where they may be recovered along with the contained zinc, copper. or lead values.
.
, ., ' ' '.'' ' ' :
PCT~CA90/00442 2~76~25 Zinc ferrite is a common component in zinc plant residues and this mineral ls partially dissolved in spent electrolyte returned from zinc electrowinning in the first step of the present process, as outllned in reaction l).
Addltional sulphuric acid may be required to maintain a solution concentration of 50 g/L ~ S04 or greater. This hot acid leach step is conveniently carried out in a series of agitated tanks at atmospheric pressure .and at a temperature of about 90O C.
ZnFe204 + 4~ SO4 , ZnS04 + Fe2 ISO4)l ~ 4~ 0 l) ~ errlc iron extracted into the hot acid leach solution is effective at leaching sulphide minerals present in the plant residue, and especially covellite, CuS. It is important that copper sulphide is leached in this step of thc process so as not to report to the zinc flotation concentrate. The reaction of ferric iron with covellite is given below.
Fe2 (50~, ) 3 ~ Cus ~ 2FeS0~ ~ CuS04 I S~ 2) To ensure rapid leaching kinetlcs, the acid level in the hot acid leach solution is maintained at 50 g/L
~ S0~ or greater. Acid in the hot acid leach discharge is partially neutralized in the next step with zinc sulphide concentrate to decr-ase reagent costs in the iron removal step and to increase the concentration of zinc in the solution. Ferric iron in the hot acid leach solution reacts with zinc sulphide and extracts zinc into the solution in a similar fashion to the reaction with covellite in the previous step. Oxygen is sparaed into the reaction vessel so as to regenerate ferric iron and continue the reaction. The net result is a decrease in the acid concentration and an increase in the zinc concentration of the solution. The equations for the reaction of zinc concentrate with ferric iron and oxygen are aiven below.
', : ' ' . . ", . ' - ' ' ,' ,, :, .. . . ~. . ,,. : .
~091/09146 PCT/CA90/nO4~2 207602~
ZnS ~ Fe2IS04) 3 ~ ZnS0- + 2FeS0~ + So 3) 2FeS04 ~ ~ S0~ ~ l/2 02 I F~ IS04 13 1 ~ 0 41 ZnS ~ ~ S04 + l/2 02 ~ ZnS04 + S + ~ O 51 After the acid concentration of the solution has been decreased to between about lO to 20 g/L ~ S04, it is desirable to reduce the majority of the iron in solution to the ferrous state so as to keep iron in solution and to minimize the consumption of sulphu~ dioxide in the reducing leach step. Reduction of ferric iron in the neutralization-reduction solution is accomplished by adding a portion of the zinc sulphide concentrate in the absence of oxyGen Ireaction 31. The neutralization-reduction step may be convenientlY carried out at atmospheric pressdure and at 35OC in a series of stirred tanks, with oxygen added to all but the last tank.
It is important that the conditions in the neutralization-reduction step be maintained sufficiently oxidlzing that copper in thc solution i9 not precipitated by metathesis with zinc sulphide, as giv-n in the reaction 6) be}ow. Copper precipitated in this manner would report to the zinc flotation concentrate in the following step of the process, rather than to the copper flotation concentrate. A concentration of between l and 5 gtL ferric iron in the neutralization-reduction solution i5 sufficient to prevent precipitation of copper in this step.
CuS0~ ~ ZnS CuS l ZnS04 6) Neutralization-reduction slurry is subjected to flotation for recovery of a zinc sulphide concentrate. The flotation concentrate contains unreacted zinc sulphide added during ncutralizat~on-reduction, zinc sulphide contained in the original zinc plant residue, and elemental sulphur produced as in leaching reactions 2) and 3). No flotation reagents are required since the elemental sulphur formed by reaction of zinc sulphide in the previous steps tends to remain on the surface of the zinc sulphide particles and assists in their flotation. The . . .
.
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~09l'09l46 PCT/CA90/~0~42 2~76~2~
_ 9 grade of the zlnc flotatlon concentrate will depend primarily uDon the quantity of excess zinc sulphide added durlng neutrallzation-reduction. The principal d_luent is elemental sulphur produced by reaction of sulphides with ferrlc iron. The zlnc flotation concentrate is sultable as feed to a roaster or to a zinc pressure leach plant. The sulphur content of the concentrate may be conveniently recovered as elemental sulphur if the concen~rate is treated in a zinc pressure leach plant.
Zinc flotation tailings slurry is the feed to a reducing leach step, where it is treated at about 100C
under an atmosphere of sulphur dioxide. Zinc, copper, iron and impurity elements in the zinc flotation tailings solids are largely extracted in this step, and .
lead found as plumbojarosite is converted to lead sulphate.
~lemental sulphur is added to precipitate copper from solution as copper sulphide. Iron in solution is also reduced to the ferrous state, allowing for a high solution concentration of iron at the low acld level (about 5 g/~
SO4). Thls ensures a minimum of iron contaminatlon of the zinc, copper and lead product streams. Appropriate reactions for the reducing leach are summarized below:
ZnFe2 04 - S02 + 2~ SO4 . ZnSO4 l 2FeS0~ 1 2~ 0 7) CuSO4 + So + S02 ~ 2HkO _ CuS ~ 2~ S04 8) PbFe6 ~SO4) 4 ~OH~ 12 ~ 3S02 ~ PbSO~ + 6FeSO4 = 6~ 0 9) Fez ISO4) 3 I 502 + 2HkO ~ 2FeSO4 1 2~ SO4 10) The reducing leach slurry is treated in a second flotation operation .to separate and recover a high grade copper sulphide concentrate, which is suitable as feed to a copper smelter. The reducing leach solution, containing the laraest part of the zinc in the original zinc plant residue, is then separated from the flotation tailin~ssolids. which contains the bulk of the lead.
The copper flotation tailings solids are subjected to another flotation operation to separate lead sulphate from the gangue minerals. The flotation of lead sulphate - - .. .. :.. ., , . :
,, . , , . . -~ . :
- - ' . '.,, ' , ' :, :
.
~091/091~6 PCT/CA90/~442 2076û2~i - lO-ma~ be ~ac ll~ated by adjustment of the pulp to a pH
between pH 8.5 and pH lO, preferably about pH 9.5. by addition of lime. followed by addition of a sulphidizing aaent such as sodium sulphide or calcium hydrosulphide, and a sultable collector such as potassium amyl xanthate.
Gypsum and other aangue minerals report to the flotation tailings, which is stockpiled.
The reducing leach solution contains iron and impurity elements which must be separated from zinc in the solution. Iron may be conveniently precipitated from the solution in a two stage process at about 85OC, and at atmospheric pressure. It i5 important for the settling and filtration characteristics of this precipitate that iron in the ~eed solution is in the ferrous state. The pH
of the solution is adjusted to between 3.5 and 4.0 in the first stage of the iron removal process. Limestone, lime, zinc calcine, basic zinc sulphate or zinc hydroxide sludge may be used for initlal pH ad~ustment and for neutralizing the acid produced upon hydrolysis of iron. Air or oxygen is added to oxidize iron to the ferric state. Ferric iron is rapidly hydrolyzed and precipitates from solution as ferric hydroxide. Impurity elements such as arsenic, germanium and fluoride are coprecipitated from solution with the iron. After liquid-solid separation, the solids, containing about 20% by weight iron. are stockpiled.
The first stage iron removal solution, containing between about O.l and l.0 g/L iron, is treated with additional neutralizing agent such as lime or zinc hydroxide sludge to raise the pH to about 5Ø The bulk of the remainder of the iron is precipitated in the presence of oxygen or air. After clarification, this second stage iron removal solution may be treated by conventional purification and electrowinning techniques for recovery of the contained zinc, or it may be integrated with the existing zinc plant. The second stage iron removal solids are recycled to the first stage of iron removal to recover .
~: .
- 1 - 2~76~2~
zlnc preclpltated at the hiaher pH ranae.
Preclous metals values are distributed between the three flotatlon concentrates. Since copper and lead smelters and zinc pressure leach plants are typlcally designed for concentration or recovery of precious metals, the silver and gold value~ in the original zinc plant residue may also be in large part recovered.
The process of the inventon will now be described with reference to the following non-limitive examples. The flrst example describes results for treatment of a zinc plant residue directly in a reducing leach step, according to the process outlined in U.S. Patent No. 4,572,822. The second example details results for treatment of the same resldue under the conditions of the present invention. .
.~ . .. . .. . . .
- . . ., . -:
: . :
() 91/1)91~6 PC~/CA90/01)~2 2~7 6~2~
A rc~e containin~ ~.4% Ca, 1.6'h Cu. 14.~% Fe, 1~.7% Pb, 0.01~0 A~, 7.5/0 S
as ~h~. 9.0% total sulpnut and 11 ~t. Zn (dry basis) wæ obtained fr~m ~e ~toc~eat an oporathg ~nc plan~ ~neral4gical ~ysis of ~e ma~rial demor~oed that calc~um was found pr~manly as gypsum. copper as copper sulphldes, iron a~ z~nc ~enrte and Jarosi~e~ b~d as lead sulpha~ and Ja ~. and 7inc as z~nc ferr~ zinc sulohsde and zinc ~il~ me re idue, o~talned as a cake contaln~ng ~6Y. by we~ght mo~tur~. was l~ched tn ss~phJric acid solubon In a 4 L aL~cl~ at 100-C wi~ ~n overpn#æure ot 27S
kPa sul~hw ~hxld~ and w~th ~on of 10 9 elemental ~utphur pet Itg of r~id~e (dry b~. The mohr ra~io of su4huric add ~ ~ c In ~o ~idUQ was 1.7:~. 'rhe resuns for a ~ten~on t~ne ot four houts ~) a~ gi~n In ~a lable bebw.
_ .
Solu~on A~ gJL ¦ Ex~on. %
- ~ -c0.01 ¦ 3~3 1 22 ~ I_ 4 1 c1 1 94 1 7~
Reduclng leach residue was subjected to flotation to recover copper as copper suiphide. The Dow rea~ent Z00 was aCde~ at a rate of 200 glt solids to ass~st Ihe flotation. ThB resulS ar~l gNen in the ta~le ~elow.
,,A~avsis. % ~
Fra~on Cu I ~ I ~ Cu ~ Zh Feed 2`3 15.1 5.~100 t 00 100 Cleaner Concen~rate 19.8 4.728.8 79 3 2 Scavencer Tailinos0.5 162 2.721 97 48 None of the produc~ o~tained in the test sequenca was of sufficient quality to a~low for tadle recovery of the contained metalls. Atthough 79h sx~ac~on of ~nc was achieved in he r~dudng teach, ~e leach solubon was dilute, cont~ining 2~ 4 g/L æn. Although a la~e frac~on of the copper in the zinc plamt residue was conver~ed to copper suJphide in ~e reduc~ng leach, the copper concen~ase recovered by flot~on of the reducing Je~ch rcsidue was also of low gra~e. containing 19.8~. Cu. Zinc sulphide was a major diluent, wi~ zinc repre~n~ng 28.8% of the weight of ths flo~ion concsntr~. Simila~y, atthough g7% of the lead was recolrered to the flo~tion tailings, this ~ilings was also of very low gra~e, containing 1 62~o by weight lead.
~f'O 91,'091 1h PCT/CA90/00442 - 13- 2~76~25 EXhJJlPLE 2 The stoc~iled z~nc plant rBsidue d~ed h Exarnple 1. recoYered a~ a fi~ter o ake containing ~6% by w~ight moisture, was repu4~d in spant elec~o~te and heated to 90-C
in a t~n of four con~nuous sb~od tank reacsors arRnge~ in cascade. The resu~ts for a r~n~on t~ns of 12 h and a spent add~on ra~ of ~m3 per torne of sto~piled res~ue (d~ ~s) are summarked in the tabb bdow. Although the tot~ retention ~me was 12 h, ~a~on was essenffally complete by tho third tank in the train, representing a retention ffme of 9 h. Zinc ex~action was limited to 34% In Shis step owins to the pre~enCB of zinc sulph~e. zinc 5;l;G~ and other zinc CO~T pounds refractory to the hot acid leach.
Solu~on Analysis, g~L ¦ Extrac~on, %
cu ! h_ I H2so~ C~ ~
32 I t 8.g ¦ 48.4 1 68.~; 1 6g ~ A
Slurry from the hot acid leach step descnbed a~ove was contacted w~h zinc sulphlde conoentrate in a traln cf fi~ agitated tanks arranged in cascaae. Oxys~en was sparged into the first four Sanks of the train. The results for addiSion of 390 kg ot zislc conc~ntrate per tonne ot stocltpiled residue ~dry basis) wlth a total retenbon time of 6 h at 90 C are sumrnarized in the table below.
Solu~on Analysis, glL
Cu ¦ Fe2+ ¦_ h ¦ H2SO~
3.4 __ ¦ 14.9 1 21.4 1 18.4 1 8A6 Sîurry from the neutrallzation-reduction step ~escrihed above w~s su~lected dlrectty to flotation. without the aid of reagen~¢. Resutts for the ffcta~on are summanzed in the table below.
Fracbon Cu ~ ~ ¦Cu ¦Pb Feed 0.7 1 3.3 16.4 100 1 00 100 Cleaner Conoent~ate 0.9 12 44.0 47.5 1.7 73.7 Scavenger Tallings 0.6 16.1 5.9 52.~ 98.3 26.3 Althou~h the zinc flotafion concentrate contained nearly 1OtD CU by weight, the quanffly of copper in ~is frac~on was leæ than that in the concentrate added to ~e neutralizaffon-reducffon step. and the zinc ~otabon concsntr~te a1d not represent an ou~et for copper contained in ~e ~nc plant residue.
, - .
-' '.
~ .
~ 0 91 /O9 1 ~h pc~/cA9o/on442 2~7~25 Zinc flotalion tallings slurry was treated at 100C, under 1C0 kPa sulohur dioxide overpressure in the re~ucin~ leacn step. Elernental sulphur was added as a Tine powder to the feed slurr~r at a rate of between 5 and 10 kg per ton of stockpiled residue. The rrtsult~ll for a retention time of 4 h are given in the ta~le below. The extraction values quoted are based on the reducing leach residue analyses. They are cumulative for the leaching steps, and metal tractions reporting to the zinc tlotation concentrate arn also included in the net 'extraction~.
_ .. _ Solids Analysis, % Extraction, %
Cu ¦ Fe ¦ Pb ¦ Zn~ Cu ¦ h ¦ Zn 2.8 1 2.3 1 21.1 ¦ 1.8c1 ¦ 89.7 1 89.8 Copper sulphide preciDitated in the reaucing leach was recovered by flotation of~he reducing leacn resibue. The ~ow reagens 7~00 was added at a rate of 200 g/t solids to assist the flotaSon. The results are given in the ta~le wi~i,ch fol 1 aws:
~nalysis, ~ = I Distribution. %
Fraction Cu Pb Zn ~ Pb i Zn feed 2.821,1 1.8 100 100 100 Cleaner Concentrate 49.0 02 1.1 86.4 0.1 3.0 Scavenaer Talllnas 0,421.5 1.8 13.6 99.9 97.0 Copper tlotation tailings was thlc~ened and the underflow filtered, washed and repulped with water to 12% solids by weight in preparation for the lead flotation step.
Slaked lime slurry wæ added to adjust the solution to pH 9 or greater, followea by sodium sulphide to activate leaC sulohate. and potassium amyl xant!nate as the collector. The results for addition of 65 kg/t lime (pH 9.8), 660 glt sodium sulphide and 800 g/t potassium amyl xanthate, with 3 minutes conditloning time between additions of reagents, are summareed in the table below- All reagent additlon rates are based on the weisht of copper flotation tailings solids.
I Analv ~ Distribution. %
Fraction ¦ Ca Cu Pb ~; Ca ¦ Cu ¦ Pb ¦ Zn Fsed 13.7 0.522.5 1.7100 100 100 1 100 Cleaner Concentrate 0.4 1.152.g 32 1.2 90.094.4 75.6 Scavenaer Tailinas 22.7 0.1 2.1 0.598.8 10.05.6 24.4 Traa~ark - .
.
.
~091/0~ PCT/CA90/00~2 2~7~
Audic zinc sulDnate solution separatçl~d trom the copper flot;~tlon tailinas solids was neutraliz~d with zinc hydroxlae sludge, obtainea by treatment of wasn solutions with lime.
under oxygen spar~lng conditions at 85~C in two stages of atmosDneric iron removal.
Sufflcisnt sludge was aaded in the first stage to maintain the soiution at pH 3.5. The first stage slurr~r wa5 filterea and the solution treated in the second stage at pH 5 by addltion of lime slurry. rhe reten~on times in these batch tests were 4 h in the first stage and 2 h in the second stage. rhQ results are given in the tabte below. Iron in the feed solution t~ the iron removal step was virtually all in the ferrous state.
Solution Al lalvsis. a/l Solids Analvsis. %
As h H2SO4 Zn.Ca ~ Zn Head 2.~9 38.2 7.5 82 . .
First Stage~0.02 0.1 pH 3.6 104 15.~ 22.5 1.1 Second Staae<0.001 <0.0005 DH 5.1 1 14. 9.9 48.4 Overall recoveries of coppe-, lead, silver and zinc in the stoc~Diled resldue to the produc~ descri~ed Exarnple 2 are summanzed in the ta~le helow:
~:E
Product G~ Ag ¦
anc flotation Concentrate ~ 2 74 16 Copper Flotation Concentrate 86 ~1 11 ~1 Lead Fhtation Concsn~ate 12 93 10 8 Low Iron anc Sulphate Solution <1 . 70 Lead Flotation Tailinas (Gypsum Residue) 1 5 5 2 Iron Piecioitate <1 4 It will be understood that changes and modifications may be made in the e.~bodiments of the invention withou~ depart-ing from the scope and purview of the appended claims - , .. . ,. .... : ....
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- I - PCT/C.~90/00442 RECOVERY OF METAL VALUES FROM Z INC PLANT RESIDUES
~IELD 0~ T~E I~ENTION
This lnventlon ~elates to the recovery of metal values from z~nc plant r~sidues and, more particularly, relates to the separation of zinc, lead. copper and precious metal values from zinc plant residues in a form amenable to the recovery of these metal values.
BACXGROUND OF THE INVENTION
Residues produced in the treatment of zinc concentrates by conventional roast-leach-electrowinning processes may contain significant ~uantlties of zinc, but the recovery of this zine may be hampered by the presence of interfering elemen~s such a~ iron or undesirable impurities such as arsenic, aermanium or fluoride. It has been well known for many years that trea~ment of zlnc ferrite residues in a hot acid leach will solubilize the zinc. Howe~er. treatment of these residues to recover the contained zinc was not economic until the 1960's when the jarosite and goethite proeesses w-re developed for handllng the iron dissolved fro~ the ferrite residues. In certain plants, the inplementatlon of these processes has allowed ~or treatment of stockpiled residues which are fed to the hot acid leach at a slow rate along with current material.
Not all zinc process residues are amenable to such a treatment, however, owing to the nature of the zinc species in the residue, which may be refractory to the hot acid leach process, or to the presence of other elements, wh~ch either interfere with the recovery of zinc or must be recovered in addition to the zinc. For exacple, the zinc plant residue may contain significant quantities of metal values other than zinc, such as lead, copper or precious metals, but the low grade of the residue in respect of these elements may preclude their economic recovery on an individual basis.
U.S. Patent No. 4,572,822 issued February 25, 1986 to Abe and Tanaka discloses a process for recovering copper from industrial residues involving a reducing leach of the residue in sulphur dioxide atmosphere to dissolve oxidtc - '' ' : .
P~T/CA90~0442 2~76025 - 2 cop~er compounds and the addition of elemental sulphur to preclpitate copper from the solutlon as cop~er sulphide.
Althouah this process allows for the recovery of copper found in the industrial residue in both sulphidic and oxidic forms as a single copper sulphide product, the process does not allow for the presence of zinc as zinc sulphide in the industrial residue. Zinc sulphlde, if present, would report along with the copper sulphide in a subsequent step to separate the copper sulphide from the industrial residue and would render the copper sulphide product of little or no value if the zinc sulphide were present in an amount equal to or greater than the amount of the copper sulphide. '!
In accordance with the process of the present lnvention, copper sulphide product from an industrial residue is recoverable substantially free of zinc.
U.S. ~atent No. 4,676,828 ~ssued June 30, 1987 to Andre discloses a process for the extraction of zinc from zinc sulphide concentrate including leaching with a dilute aqueous solution of sulphuric acid under atmospheric pressure. This process, however, relates to the extractlon of zinc and copper from sulphurous zinciferous materials, and does not permit the separation and recovery of copper and zinc from industrial residues which contain a large proportion of oxidic compounds.
The flotation of lead sulphate and the precipitation of iron from a zinc sulphate solution are discussed in Papers by Fuerstenau et al, ''The Surface Characteristics and Flotation Behavior of Anglesite and Cerussite", Int. J. of Min. Proc., 1987, 20, 73-85: Andre and Masson, "The Goethite Process in Retreating Zinc Leaching Residues", presented at the 102nd AIME Annual Meeting, 1973, Chicago; 30xall and James, "Experience with the Goethite Process at National Zinc"; and in J.E.
Dutrizac and A.J. Monhemius ~Ed.), Iron Control in Hydrometallurgy, 1986, Ellis Horwood, Chichester, 676-686.
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: .
~O 91/091 ~h PCT/CA90/0 ~ 2 ~ 3 ~ 20~25 None of these Papers discloses or suagesls the comblnatlon of steps and parameters of the process c' the present lnvention for the recovery of metal values from industrial residues The process of the invention permits separation of zinc. copper, lead and precious metals from zinc plant resldues in a readily recoverable form and surprisingly allows for recovery of metal values present in the head material in two or more substantially different mineral forms. Zinc or copper, for example. may be recovered in high yields whether they are present as sulphidic or oxidic compounds, or both.
SUMMARY OF T~E INVENTION
In accordance with the process of the present inventior.. -inc plant residue is treated in both hot acid leaching and reducing leaching stages, with a neutralization and reduction stage and a flotation stage between the two leaching stages. Copper sulphide in the residue is first reacted with ferric iron in solution in the hot acid leach stage to oxidize the copper sulphide and to brina the copper conte~t of this sulphide into the solution, Zinc sulphide in the residue is then recovered by flotation separation prio- to leaching the remaining oxidic copper compounds and precipitating copper from the solution in the reducing leach stage with sulphur dioxide and elemental sulphur. The copper sulphide product that is recovered is substantially free of zinc. As a result of this combination of steps, both copper and zinc may be present in oxidic and sulphidic forms in the feed to the process and be recovered separately to hi~h grade concentrates or to a high stren~th solution.
More particularly, the process of the invention relates to the recovery of zinc, lead, copper and precious metals from zinc plant residue containing ferrites compris-ina the steps of leaching said zinc plant residues with return spent electrolyte containina ~ S04 in an amount effective to dissolve the ferrites and to maintain at least 50 g/L ~ S04 in a hot acid leach z: a temperature in the range of . . ~ ,. ' , .
.~ ~ .. . .
., .
~ ' .
,.:
~ 0 91/09146 PCT/CA90/1)0~142 2 ~ 7 6 ~ 2 ~
70O to 100C and at atmospherlc pressure to partlally dissolve zinc, copper. iron and impurlty elements and to e~sentially leach sulphide copper; treating the resulting 10ach slurry with zinc concentrate under oxidizing ccnditions at atmospheric pressure at a temperature in the range of 70 to 100C to consume excess acid from said hot acid leach and to increase the concentration of zinc in the leach solution and continulng said treatment with zinc concentrate under reducing conditions to reduce ferric iron in solution to ferrous iron; recovering excess zinc concentrate and elemental sulphur produced in the reaction of zinc concentrate with acid and ferric iron by zinc flotation as a flotation concentrate: subjecting ~inc flotation tailings to a reducing leach in the presence of gaseous sulphur dioxide and elemental sulphur at a temperature in the range of 700 to 120C with a sulphur dioxide overpressure of at least 30 kPa to extract zinc, copper, iron and impurity elements, to reprecipitate copper as coppor sulphlde, and to convert load in jarosite to lead sulphate: recovering copper sulphlde by copper flotation as a ~lotation concentrate: subjecting the copper flotation tailings to a liguid-solid separation;
reco~ering lead sulphate as a flotation concentrate by lead flotation of the separated solids from the liquid-solid separation of the copper flotation tailings; and recovering and treating the reducing leach solution from the liquid-solid separation for the recover~ of contained zinc values.
The process preferably includes adjusting the pH of the reducing leach solution from the liquid-solid separation to about 3.5 to 4.0 and oxidizing the largest part of contained ferrous iron to ferric iron in an oxidizing atmosphere at a temperature in the range of 700 to 100C. preferably about 85oC, at atmospheric pressure for precipitation of iron as ferric hydroxide or hydrated ferric oxide with impurity elements, and separating the " : ' ' '' - ~ ' ~ ' ., ' ; ' , ' -, ~091/09116 PCT/CA90/00~42 - 5 _ 2976~2~
resldual solutlon: and adjustin~ the pH of the recovered reSldUal sOlUtlOn tO about 5.0, oxidizing the contained iron to ferric iron in an oxidizing atmosDhere at a temperature in the range of 700 ro 100C, preferably about 8~oC. at atmospheric Dressure for precipitation of iron as ferric hydroxide or hydrated ferric oxide, and separating the residual solution for recovery of contained zinc values.
DESCRIPTION OF THE ~RAWING
The process of the invention will be described with reference to the flow sheet of the accompanying drawing.
DESCRIPTION OF THE PREFERRED EMBODIMENT
Zinc plant residue containing zinc, lead, copper and precious metal values is contacted with return zinc spent electrolyte at atmospheric pressure and a temperature of at least 70C, preferably in the range of 700 to 100C, and more preferably, 850 to 950C to partially dissolve zinc, copper, iron and impurity elements. It is of particular importance to the recovery of copper that sulphide copper in the zinc pla~t residue is leached ln this stcp, The resultins slurry i9 treated with zinc concentrate under oxygen sparging conditions at atmospheric pressure and a temperature of at least 700C or more preferably 850 to 950C, to consume the excess of acid from the previous step and to increase the concentration of zinc in the solution. Additional zinc concentrate is fed to the slurry in the absence of oxygen, but at otherwise similar temperature and pressure, to reduce ferric iron in solution to the ferrous state. Excess zinc concentrate and elemental sulphur produced in .the reaction of zinc concentrate with acid and ferric iron is recovered by flotation, this flotation concentrate being suitable as feed to a zinc pressure leach autoclave or a roaster.
The flotation tailings is treated with sulphur dioxide and elemental sulphur at a temperature of at least 700C. preferablY in the range of 70 to 120C, and more .. :, ' ' ' :. . , .. ,. - ~, . .
' ' '"'" ' . ' :... ' ~, . , :
~091/09~6 PCT/CA90/n~2 207~02~ - 6 -~rererablv ln t~e ranae sOo ~o 100C and wlth an overpressure of sulphur dioxide of at least 30 kPa.
pre~erably at about 100 kPa. to further extract zlnc.
copper. lron and impurity elements. to repreclpitate copper as copper sulphide. and to convert lead in jarosite to lead sulphate. The slurry resulting from this reducina leach is sub~ected to a second flotation step to separate a copper sulphide concentrate, this flotation concentrate being suitable as feed to a copper smelter. After liq~id-solid separation and adjustment to between pH 8.5 and p~ 10.0, more prefrably pH 9.S, the flotation tailings solids are sulphidized for recovery of a lead sulphate flotatlon concentrate. which is suitable as feed to a lead smelter.
The lead flotation tailings, comprised of aypsum and other gangue minerals, is stoc~piled.
The leach solution separated from the copper flotation tailings solids, containing the largest part of the zinc in the head material, is treated with lime or zinc hydroxide sludge in a two stage process for precipitation of iron and impurity elements from the s~lution. Both stages are maintained at a temperature between 70 and 100C, more preferably at 85oC. The solution is maintained at about pH 4 or less in the first stage to limit the coprecipitation of zinc. After liquid-solid separation, the iron cake is impounded and the solution sent to a second stage of iron removal at pH 5. The resulting slurry is clarified and the solids recycled to the first stage of the iron removal step for recovery of precipitated zinc.
Zinc may be recoverd from the low iron zinc sulphate solution using conventional purification and electrowinning techniques, or the solution may be integrated with the existing zinc plant. Precious metal values in the oxide residue head material are distributed among the three flotation concentrates in the process, where they may be recovered along with the contained zinc, copper. or lead values.
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, ., ' ' '.'' ' ' :
PCT~CA90/00442 2~76~25 Zinc ferrite is a common component in zinc plant residues and this mineral ls partially dissolved in spent electrolyte returned from zinc electrowinning in the first step of the present process, as outllned in reaction l).
Addltional sulphuric acid may be required to maintain a solution concentration of 50 g/L ~ S04 or greater. This hot acid leach step is conveniently carried out in a series of agitated tanks at atmospheric pressure .and at a temperature of about 90O C.
ZnFe204 + 4~ SO4 , ZnS04 + Fe2 ISO4)l ~ 4~ 0 l) ~ errlc iron extracted into the hot acid leach solution is effective at leaching sulphide minerals present in the plant residue, and especially covellite, CuS. It is important that copper sulphide is leached in this step of thc process so as not to report to the zinc flotation concentrate. The reaction of ferric iron with covellite is given below.
Fe2 (50~, ) 3 ~ Cus ~ 2FeS0~ ~ CuS04 I S~ 2) To ensure rapid leaching kinetlcs, the acid level in the hot acid leach solution is maintained at 50 g/L
~ S0~ or greater. Acid in the hot acid leach discharge is partially neutralized in the next step with zinc sulphide concentrate to decr-ase reagent costs in the iron removal step and to increase the concentration of zinc in the solution. Ferric iron in the hot acid leach solution reacts with zinc sulphide and extracts zinc into the solution in a similar fashion to the reaction with covellite in the previous step. Oxygen is sparaed into the reaction vessel so as to regenerate ferric iron and continue the reaction. The net result is a decrease in the acid concentration and an increase in the zinc concentration of the solution. The equations for the reaction of zinc concentrate with ferric iron and oxygen are aiven below.
', : ' ' . . ", . ' - ' ' ,' ,, :, .. . . ~. . ,,. : .
~091/09146 PCT/CA90/nO4~2 207602~
ZnS ~ Fe2IS04) 3 ~ ZnS0- + 2FeS0~ + So 3) 2FeS04 ~ ~ S0~ ~ l/2 02 I F~ IS04 13 1 ~ 0 41 ZnS ~ ~ S04 + l/2 02 ~ ZnS04 + S + ~ O 51 After the acid concentration of the solution has been decreased to between about lO to 20 g/L ~ S04, it is desirable to reduce the majority of the iron in solution to the ferrous state so as to keep iron in solution and to minimize the consumption of sulphu~ dioxide in the reducing leach step. Reduction of ferric iron in the neutralization-reduction solution is accomplished by adding a portion of the zinc sulphide concentrate in the absence of oxyGen Ireaction 31. The neutralization-reduction step may be convenientlY carried out at atmospheric pressdure and at 35OC in a series of stirred tanks, with oxygen added to all but the last tank.
It is important that the conditions in the neutralization-reduction step be maintained sufficiently oxidlzing that copper in thc solution i9 not precipitated by metathesis with zinc sulphide, as giv-n in the reaction 6) be}ow. Copper precipitated in this manner would report to the zinc flotation concentrate in the following step of the process, rather than to the copper flotation concentrate. A concentration of between l and 5 gtL ferric iron in the neutralization-reduction solution i5 sufficient to prevent precipitation of copper in this step.
CuS0~ ~ ZnS CuS l ZnS04 6) Neutralization-reduction slurry is subjected to flotation for recovery of a zinc sulphide concentrate. The flotation concentrate contains unreacted zinc sulphide added during ncutralizat~on-reduction, zinc sulphide contained in the original zinc plant residue, and elemental sulphur produced as in leaching reactions 2) and 3). No flotation reagents are required since the elemental sulphur formed by reaction of zinc sulphide in the previous steps tends to remain on the surface of the zinc sulphide particles and assists in their flotation. The . . .
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~09l'09l46 PCT/CA90/~0~42 2~76~2~
_ 9 grade of the zlnc flotatlon concentrate will depend primarily uDon the quantity of excess zinc sulphide added durlng neutrallzation-reduction. The principal d_luent is elemental sulphur produced by reaction of sulphides with ferrlc iron. The zlnc flotation concentrate is sultable as feed to a roaster or to a zinc pressure leach plant. The sulphur content of the concentrate may be conveniently recovered as elemental sulphur if the concen~rate is treated in a zinc pressure leach plant.
Zinc flotation tailings slurry is the feed to a reducing leach step, where it is treated at about 100C
under an atmosphere of sulphur dioxide. Zinc, copper, iron and impurity elements in the zinc flotation tailings solids are largely extracted in this step, and .
lead found as plumbojarosite is converted to lead sulphate.
~lemental sulphur is added to precipitate copper from solution as copper sulphide. Iron in solution is also reduced to the ferrous state, allowing for a high solution concentration of iron at the low acld level (about 5 g/~
SO4). Thls ensures a minimum of iron contaminatlon of the zinc, copper and lead product streams. Appropriate reactions for the reducing leach are summarized below:
ZnFe2 04 - S02 + 2~ SO4 . ZnSO4 l 2FeS0~ 1 2~ 0 7) CuSO4 + So + S02 ~ 2HkO _ CuS ~ 2~ S04 8) PbFe6 ~SO4) 4 ~OH~ 12 ~ 3S02 ~ PbSO~ + 6FeSO4 = 6~ 0 9) Fez ISO4) 3 I 502 + 2HkO ~ 2FeSO4 1 2~ SO4 10) The reducing leach slurry is treated in a second flotation operation .to separate and recover a high grade copper sulphide concentrate, which is suitable as feed to a copper smelter. The reducing leach solution, containing the laraest part of the zinc in the original zinc plant residue, is then separated from the flotation tailin~ssolids. which contains the bulk of the lead.
The copper flotation tailings solids are subjected to another flotation operation to separate lead sulphate from the gangue minerals. The flotation of lead sulphate - - .. .. :.. ., , . :
,, . , , . . -~ . :
- - ' . '.,, ' , ' :, :
.
~091/091~6 PCT/CA90/~442 2076û2~i - lO-ma~ be ~ac ll~ated by adjustment of the pulp to a pH
between pH 8.5 and pH lO, preferably about pH 9.5. by addition of lime. followed by addition of a sulphidizing aaent such as sodium sulphide or calcium hydrosulphide, and a sultable collector such as potassium amyl xanthate.
Gypsum and other aangue minerals report to the flotation tailings, which is stockpiled.
The reducing leach solution contains iron and impurity elements which must be separated from zinc in the solution. Iron may be conveniently precipitated from the solution in a two stage process at about 85OC, and at atmospheric pressure. It i5 important for the settling and filtration characteristics of this precipitate that iron in the ~eed solution is in the ferrous state. The pH
of the solution is adjusted to between 3.5 and 4.0 in the first stage of the iron removal process. Limestone, lime, zinc calcine, basic zinc sulphate or zinc hydroxide sludge may be used for initlal pH ad~ustment and for neutralizing the acid produced upon hydrolysis of iron. Air or oxygen is added to oxidize iron to the ferric state. Ferric iron is rapidly hydrolyzed and precipitates from solution as ferric hydroxide. Impurity elements such as arsenic, germanium and fluoride are coprecipitated from solution with the iron. After liquid-solid separation, the solids, containing about 20% by weight iron. are stockpiled.
The first stage iron removal solution, containing between about O.l and l.0 g/L iron, is treated with additional neutralizing agent such as lime or zinc hydroxide sludge to raise the pH to about 5Ø The bulk of the remainder of the iron is precipitated in the presence of oxygen or air. After clarification, this second stage iron removal solution may be treated by conventional purification and electrowinning techniques for recovery of the contained zinc, or it may be integrated with the existing zinc plant. The second stage iron removal solids are recycled to the first stage of iron removal to recover .
~: .
- 1 - 2~76~2~
zlnc preclpltated at the hiaher pH ranae.
Preclous metals values are distributed between the three flotatlon concentrates. Since copper and lead smelters and zinc pressure leach plants are typlcally designed for concentration or recovery of precious metals, the silver and gold value~ in the original zinc plant residue may also be in large part recovered.
The process of the inventon will now be described with reference to the following non-limitive examples. The flrst example describes results for treatment of a zinc plant residue directly in a reducing leach step, according to the process outlined in U.S. Patent No. 4,572,822. The second example details results for treatment of the same resldue under the conditions of the present invention. .
.~ . .. . .. . . .
- . . ., . -:
: . :
() 91/1)91~6 PC~/CA90/01)~2 2~7 6~2~
A rc~e containin~ ~.4% Ca, 1.6'h Cu. 14.~% Fe, 1~.7% Pb, 0.01~0 A~, 7.5/0 S
as ~h~. 9.0% total sulpnut and 11 ~t. Zn (dry basis) wæ obtained fr~m ~e ~toc~eat an oporathg ~nc plan~ ~neral4gical ~ysis of ~e ma~rial demor~oed that calc~um was found pr~manly as gypsum. copper as copper sulphldes, iron a~ z~nc ~enrte and Jarosi~e~ b~d as lead sulpha~ and Ja ~. and 7inc as z~nc ferr~ zinc sulohsde and zinc ~il~ me re idue, o~talned as a cake contaln~ng ~6Y. by we~ght mo~tur~. was l~ched tn ss~phJric acid solubon In a 4 L aL~cl~ at 100-C wi~ ~n overpn#æure ot 27S
kPa sul~hw ~hxld~ and w~th ~on of 10 9 elemental ~utphur pet Itg of r~id~e (dry b~. The mohr ra~io of su4huric add ~ ~ c In ~o ~idUQ was 1.7:~. 'rhe resuns for a ~ten~on t~ne ot four houts ~) a~ gi~n In ~a lable bebw.
_ .
Solu~on A~ gJL ¦ Ex~on. %
- ~ -c0.01 ¦ 3~3 1 22 ~ I_ 4 1 c1 1 94 1 7~
Reduclng leach residue was subjected to flotation to recover copper as copper suiphide. The Dow rea~ent Z00 was aCde~ at a rate of 200 glt solids to ass~st Ihe flotation. ThB resulS ar~l gNen in the ta~le ~elow.
,,A~avsis. % ~
Fra~on Cu I ~ I ~ Cu ~ Zh Feed 2`3 15.1 5.~100 t 00 100 Cleaner Concen~rate 19.8 4.728.8 79 3 2 Scavencer Tailinos0.5 162 2.721 97 48 None of the produc~ o~tained in the test sequenca was of sufficient quality to a~low for tadle recovery of the contained metalls. Atthough 79h sx~ac~on of ~nc was achieved in he r~dudng teach, ~e leach solubon was dilute, cont~ining 2~ 4 g/L æn. Although a la~e frac~on of the copper in the zinc plamt residue was conver~ed to copper suJphide in ~e reduc~ng leach, the copper concen~ase recovered by flot~on of the reducing Je~ch rcsidue was also of low gra~e. containing 19.8~. Cu. Zinc sulphide was a major diluent, wi~ zinc repre~n~ng 28.8% of the weight of ths flo~ion concsntr~. Simila~y, atthough g7% of the lead was recolrered to the flo~tion tailings, this ~ilings was also of very low gra~e, containing 1 62~o by weight lead.
~f'O 91,'091 1h PCT/CA90/00442 - 13- 2~76~25 EXhJJlPLE 2 The stoc~iled z~nc plant rBsidue d~ed h Exarnple 1. recoYered a~ a fi~ter o ake containing ~6% by w~ight moisture, was repu4~d in spant elec~o~te and heated to 90-C
in a t~n of four con~nuous sb~od tank reacsors arRnge~ in cascade. The resu~ts for a r~n~on t~ns of 12 h and a spent add~on ra~ of ~m3 per torne of sto~piled res~ue (d~ ~s) are summarked in the tabb bdow. Although the tot~ retention ~me was 12 h, ~a~on was essenffally complete by tho third tank in the train, representing a retention ffme of 9 h. Zinc ex~action was limited to 34% In Shis step owins to the pre~enCB of zinc sulph~e. zinc 5;l;G~ and other zinc CO~T pounds refractory to the hot acid leach.
Solu~on Analysis, g~L ¦ Extrac~on, %
cu ! h_ I H2so~ C~ ~
32 I t 8.g ¦ 48.4 1 68.~; 1 6g ~ A
Slurry from the hot acid leach step descnbed a~ove was contacted w~h zinc sulphlde conoentrate in a traln cf fi~ agitated tanks arranged in cascaae. Oxys~en was sparged into the first four Sanks of the train. The results for addiSion of 390 kg ot zislc conc~ntrate per tonne ot stocltpiled residue ~dry basis) wlth a total retenbon time of 6 h at 90 C are sumrnarized in the table below.
Solu~on Analysis, glL
Cu ¦ Fe2+ ¦_ h ¦ H2SO~
3.4 __ ¦ 14.9 1 21.4 1 18.4 1 8A6 Sîurry from the neutrallzation-reduction step ~escrihed above w~s su~lected dlrectty to flotation. without the aid of reagen~¢. Resutts for the ffcta~on are summanzed in the table below.
Fracbon Cu ~ ~ ¦Cu ¦Pb Feed 0.7 1 3.3 16.4 100 1 00 100 Cleaner Conoent~ate 0.9 12 44.0 47.5 1.7 73.7 Scavenger Tallings 0.6 16.1 5.9 52.~ 98.3 26.3 Althou~h the zinc flotafion concentrate contained nearly 1OtD CU by weight, the quanffly of copper in ~is frac~on was leæ than that in the concentrate added to ~e neutralizaffon-reducffon step. and the zinc ~otabon concsntr~te a1d not represent an ou~et for copper contained in ~e ~nc plant residue.
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-' '.
~ .
~ 0 91 /O9 1 ~h pc~/cA9o/on442 2~7~25 Zinc flotalion tallings slurry was treated at 100C, under 1C0 kPa sulohur dioxide overpressure in the re~ucin~ leacn step. Elernental sulphur was added as a Tine powder to the feed slurr~r at a rate of between 5 and 10 kg per ton of stockpiled residue. The rrtsult~ll for a retention time of 4 h are given in the ta~le below. The extraction values quoted are based on the reducing leach residue analyses. They are cumulative for the leaching steps, and metal tractions reporting to the zinc tlotation concentrate arn also included in the net 'extraction~.
_ .. _ Solids Analysis, % Extraction, %
Cu ¦ Fe ¦ Pb ¦ Zn~ Cu ¦ h ¦ Zn 2.8 1 2.3 1 21.1 ¦ 1.8c1 ¦ 89.7 1 89.8 Copper sulphide preciDitated in the reaucing leach was recovered by flotation of~he reducing leacn resibue. The ~ow reagens 7~00 was added at a rate of 200 g/t solids to assist the flotaSon. The results are given in the ta~le wi~i,ch fol 1 aws:
~nalysis, ~ = I Distribution. %
Fraction Cu Pb Zn ~ Pb i Zn feed 2.821,1 1.8 100 100 100 Cleaner Concentrate 49.0 02 1.1 86.4 0.1 3.0 Scavenaer Talllnas 0,421.5 1.8 13.6 99.9 97.0 Copper tlotation tailings was thlc~ened and the underflow filtered, washed and repulped with water to 12% solids by weight in preparation for the lead flotation step.
Slaked lime slurry wæ added to adjust the solution to pH 9 or greater, followea by sodium sulphide to activate leaC sulohate. and potassium amyl xant!nate as the collector. The results for addition of 65 kg/t lime (pH 9.8), 660 glt sodium sulphide and 800 g/t potassium amyl xanthate, with 3 minutes conditloning time between additions of reagents, are summareed in the table below- All reagent additlon rates are based on the weisht of copper flotation tailings solids.
I Analv ~ Distribution. %
Fraction ¦ Ca Cu Pb ~; Ca ¦ Cu ¦ Pb ¦ Zn Fsed 13.7 0.522.5 1.7100 100 100 1 100 Cleaner Concentrate 0.4 1.152.g 32 1.2 90.094.4 75.6 Scavenaer Tailinas 22.7 0.1 2.1 0.598.8 10.05.6 24.4 Traa~ark - .
.
.
~091/0~ PCT/CA90/00~2 2~7~
Audic zinc sulDnate solution separatçl~d trom the copper flot;~tlon tailinas solids was neutraliz~d with zinc hydroxlae sludge, obtainea by treatment of wasn solutions with lime.
under oxygen spar~lng conditions at 85~C in two stages of atmosDneric iron removal.
Sufflcisnt sludge was aaded in the first stage to maintain the soiution at pH 3.5. The first stage slurr~r wa5 filterea and the solution treated in the second stage at pH 5 by addltion of lime slurry. rhe reten~on times in these batch tests were 4 h in the first stage and 2 h in the second stage. rhQ results are given in the tabte below. Iron in the feed solution t~ the iron removal step was virtually all in the ferrous state.
Solution Al lalvsis. a/l Solids Analvsis. %
As h H2SO4 Zn.Ca ~ Zn Head 2.~9 38.2 7.5 82 . .
First Stage~0.02 0.1 pH 3.6 104 15.~ 22.5 1.1 Second Staae<0.001 <0.0005 DH 5.1 1 14. 9.9 48.4 Overall recoveries of coppe-, lead, silver and zinc in the stoc~Diled resldue to the produc~ descri~ed Exarnple 2 are summanzed in the ta~le helow:
~:E
Product G~ Ag ¦
anc flotation Concentrate ~ 2 74 16 Copper Flotation Concentrate 86 ~1 11 ~1 Lead Fhtation Concsn~ate 12 93 10 8 Low Iron anc Sulphate Solution <1 . 70 Lead Flotation Tailinas (Gypsum Residue) 1 5 5 2 Iron Piecioitate <1 4 It will be understood that changes and modifications may be made in the e.~bodiments of the invention withou~ depart-ing from the scope and purview of the appended claims - , .. . ,. .... : ....
.
.
Claims (8)
1. A process for the recovery of zinc, lead, copper and precious metals from zinc plant residues containing ferrites comprising the steps of:
leaching said zinc plant residues with return spent electrolyte containing H2 SO4 in an amount effective to dissolve the ferrites and to maintain at least 50 g/L H2 SO4 in a hot acid leach at a temperature in the range of 70° to 100°C and at atmospheric pressure to partially dissolve zinc, copper, iron and impurity elements and to essentially leach sulphide copper:
treating the resulting leach slurry with zinc concentrate under oxidizing conditions at atmospheric pressure at a temperature in the range of 70° to 100° C to consume excess acid from said hot acid leach and to increase the concentration of zinc in the leach solution and continuing said treatment with zinc concentrate under reducing conditions to reduce ferric iron in solution to ferrous iron:
recovering excess zinc concentrate and elemental sulphur produced in the reaction of zinc concentrate with acid and ferric iron by zinc flotation as a flotation concentrate;
subjecting zinc flotation tailings to a reducing leach in the presence of gaseous sulphur dioxide and elemental sulphur at a temperature in the range of 70° to 120°C with a sulphur dioxide overpressure of at least 30 kPa to extract zinc, copper, iron and impurity elements, to reprecipitate copper as copper sulphide, and to convert lead in jarosite to lead sulphate;
recovering copper sulphide by copper flotation as a flotation concentrate;
subjecting the copper flotation tailings to a liquid-solid separation;
recovering lead sulphate as a flotation concentrate by lead flotation of the separated solids from the liquid-solid separation of the copper flotation tailings; and recovering and treating the reducing leach solution from the liquid-solid separation for the recovery of contained zinc values.
leaching said zinc plant residues with return spent electrolyte containing H2 SO4 in an amount effective to dissolve the ferrites and to maintain at least 50 g/L H2 SO4 in a hot acid leach at a temperature in the range of 70° to 100°C and at atmospheric pressure to partially dissolve zinc, copper, iron and impurity elements and to essentially leach sulphide copper:
treating the resulting leach slurry with zinc concentrate under oxidizing conditions at atmospheric pressure at a temperature in the range of 70° to 100° C to consume excess acid from said hot acid leach and to increase the concentration of zinc in the leach solution and continuing said treatment with zinc concentrate under reducing conditions to reduce ferric iron in solution to ferrous iron:
recovering excess zinc concentrate and elemental sulphur produced in the reaction of zinc concentrate with acid and ferric iron by zinc flotation as a flotation concentrate;
subjecting zinc flotation tailings to a reducing leach in the presence of gaseous sulphur dioxide and elemental sulphur at a temperature in the range of 70° to 120°C with a sulphur dioxide overpressure of at least 30 kPa to extract zinc, copper, iron and impurity elements, to reprecipitate copper as copper sulphide, and to convert lead in jarosite to lead sulphate;
recovering copper sulphide by copper flotation as a flotation concentrate;
subjecting the copper flotation tailings to a liquid-solid separation;
recovering lead sulphate as a flotation concentrate by lead flotation of the separated solids from the liquid-solid separation of the copper flotation tailings; and recovering and treating the reducing leach solution from the liquid-solid separation for the recovery of contained zinc values.
2. In a process as claimed in claim 1, adjusting the pH of the reducing leach solution from the liquid-solid separation to about 3.5 to 4.0, and oxidizing the largest part of contained ferrous iron to ferric iron in an oxidizing atmosphere at a temperature of about 85°C at atmospheric pressure for precipitation of iron as ferric hydroxide or hydrated ferric oxide with impurity elements, and separating the residual solution; and adjusting the pH
of the recovered residual solution to about 5.0, oxidizing the contained iron to ferric iron in an oxidizing atmosphere at a temperature of about 85°C at atmospheric pressure for precipitation of iron as ferric hydroxide or hydrated ferric oxide, and separating the residual solution for recovery of contained zinc values.
of the recovered residual solution to about 5.0, oxidizing the contained iron to ferric iron in an oxidizing atmosphere at a temperature of about 85°C at atmospheric pressure for precipitation of iron as ferric hydroxide or hydrated ferric oxide, and separating the residual solution for recovery of contained zinc values.
3. A process as claimed in claim 1 or 2 in which said zinc plant residues are leached in a hot acid leach at a temperature in the range of 85° to 95°C.
4. A process as claimed in claim 1 or 2 in which said zinc plant residues are leached in a hot acid leach at a temperature in the range of 85° to 95°C and the leach slurry is treated with zinc concentrate at a temperature in the range of 85° to 95°C.
5. A process as claimed in claim 1 or 2 in which said zinc plant residues are leached in a hot acid leach at a temperature in the range of 85° to 95°C, leach slurry is treated with zinc concentrate at a temperature in the range of 85° to 95°C, and zinc flotation tailings are subjected to a reducing leach at a temperature at about 100°C and at a sulphur dioxide overpressure of about 100 kPa.
6. A process as claimed in claim 1 or 2 in which said impurity elements are arsenic, germanium and fluoride.
7. A process as claimed in claim 1 or 2 in which the pH is adjusted by the addition of limestone, lime, zinc calcine, basic zinc sulphate or zinc hydroxide sludge.
8. A process as claimed in claim 1 or 2 in which the oxidizing atmosphere is provided by oxygen or air.
Applications Claiming Priority (2)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| GB8928368A GB8928368D0 (en) | 1989-12-15 | 1989-12-15 | Recovery of metal values from zinc plant residues |
| GB8928368.3 | 1989-12-15 |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CA2076025A1 true CA2076025A1 (en) | 1991-06-16 |
Family
ID=10667999
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CA 2076025 Abandoned CA2076025A1 (en) | 1989-12-15 | 1990-12-14 | Recovery of metal values from zinc plant residues |
Country Status (6)
| Country | Link |
|---|---|
| US (1) | US5348713A (en) |
| AU (1) | AU6913191A (en) |
| CA (1) | CA2076025A1 (en) |
| GB (1) | GB8928368D0 (en) |
| WO (1) | WO1991009146A1 (en) |
| ZA (1) | ZA9010122B (en) |
Families Citing this family (27)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| GB9306201D0 (en) * | 1993-03-25 | 1993-05-19 | Sherritt Gordon Ltd | Recovery of zinc,iron,lead and silver values from sinc sulphide concentrate by a multi-stage pressure oxidation process |
| WO1995006140A1 (en) * | 1993-08-27 | 1995-03-02 | N.V. Union Miniere S.A. | Method for removing zinc from sulphured concentrates |
| BE1007906A3 (en) * | 1993-12-23 | 1995-11-14 | Union Miniere Sa | Operational method for leaching zinc concentrate and zinc ferritecontaining sulphur |
| BE1007417A3 (en) * | 1993-08-27 | 1995-06-13 | Union Miniere Sa | Leaching method for zinc sulphide concentrate and zinc ferrite |
| GB9615946D0 (en) * | 1996-07-30 | 1996-09-11 | Tetra Europ Ltd | Zinc recovery process |
| JP4129499B2 (en) * | 2000-08-08 | 2008-08-06 | Ｄｏｗａメタルマイン株式会社 | Method for recovering sulfur from minerals |
| US6451088B1 (en) * | 2001-07-25 | 2002-09-17 | Phelps Dodge Corporation | Method for improving metals recovery using high temperature leaching |
| CN100362691C (en) * | 2006-01-10 | 2008-01-16 | 赵恒祥 | Cyclic utilization method and apparatus of lead dioxide from storage cells |
| WO2007087675A1 (en) * | 2006-01-31 | 2007-08-09 | Murrin Murrin Operations Pty Ltd | Improved base metal recovery process from heap leaching |
| CN103397200B (en) * | 2013-08-23 | 2015-06-10 | 阳谷祥光铜业有限公司 | Method for removing lead, zinc, arsenic, antimony, bismuth and tin from copper matte |
| CN104789783B (en) * | 2015-05-18 | 2017-04-19 | 郴州市金贵银业股份有限公司 | Process for selective efficient copper extraction and comprehensive recovery from lead copper matte |
| CN105525097B (en) * | 2016-01-21 | 2017-11-10 | 昆明理工大学 | A kind of method and apparatus that zinc leaching residue is leached with sulfur dioxide controlling potential method |
| CN108330281A (en) * | 2018-03-16 | 2018-07-27 | 云锡文山锌铟冶炼有限公司 | Zinc abstraction reducing leaching device |
| CN109055749B (en) * | 2018-09-28 | 2020-06-16 | 长沙有色冶金设计研究院有限公司 | Method for treating zinc oxygen pressure leaching solution |
| CN111876612A (en) * | 2020-07-14 | 2020-11-03 | 矿冶科技集团有限公司 | Method for treating zinc-iron-containing acidic solution |
| CN112501452B (en) * | 2020-11-05 | 2022-08-19 | 矿冶科技集团有限公司 | Method for purifying zinc mineral material leaching solution |
| CN112934474B (en) * | 2021-03-08 | 2023-11-07 | 江苏北矿金属循环利用科技有限公司 | Method for recycling sulfur by flocculation flotation of zinc leached high-sulfur slag |
| CN113546754B (en) * | 2021-07-23 | 2022-10-11 | 昆明冶金研究院有限公司 | Process for comprehensively utilizing oxygen-sulfur mixed lead-zinc ore |
| CN113832346B (en) * | 2021-09-16 | 2023-07-21 | 云南驰宏资源综合利用有限公司 | Method for efficiently and simply treating germanium-containing zinc leaching residues |
| CN113897491B (en) * | 2021-09-16 | 2023-05-02 | 昆明理工大学 | Method for comprehensively and efficiently treating zinc leaching slag |
| CN114214520B (en) * | 2021-12-17 | 2024-04-16 | 蒙自矿冶有限责任公司 | Copper-containing refractory material waste-free environment-friendly recovery method |
| CN114438318B (en) * | 2021-12-30 | 2023-12-08 | 云锡文山锌铟冶炼有限公司 | Zinc hydrometallurgy start-up method |
| CN114438328B (en) * | 2021-12-30 | 2023-09-22 | 云锡文山锌铟冶炼有限公司 | Device and method for producing zinc sulfate leaching solution in zinc hydrometallurgy process |
| CN114561542B (en) * | 2022-02-25 | 2024-02-02 | 盛隆资源再生（无锡）有限公司 | Method for improving copper recovery rate in iron-containing copper waste liquid |
| CN114671411B (en) * | 2022-04-18 | 2023-07-28 | 中南大学 | Method for separating heavy metal and sulfur in sulfur-oxygen-containing pressure leaching slag |
| CN115874058B (en) * | 2023-03-02 | 2023-05-12 | 昆明理工大学 | Method for efficiently enriching germanium by pre-dezincification and germanium-containing zinc oxide smoke neutralization method |
| CN116812874B (en) * | 2023-08-30 | 2023-11-17 | 昆明理工大学 | Method for efficiently recycling sulfur and zinc and silver from zinc hydrometallurgy high-sulfur residues |
Family Cites Families (11)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US3510414A (en) * | 1967-12-04 | 1970-05-05 | Bunker Hill Co | Process of recovering metals from zinc plant residue |
| IE32587B1 (en) * | 1968-11-20 | 1973-09-19 | Mines Fond Zinc Vieille | Improvements in or relating to valorization |
| BE756611A (en) * | 1969-09-25 | 1971-03-01 | Ruhr Zink Gmbh | PROCESS FOR OBTAINING ZINC |
| GB1509537A (en) * | 1974-09-13 | 1978-05-04 | Cominco Ltd | Treatment of zinc plant residues |
| US4149945A (en) * | 1978-04-13 | 1979-04-17 | Kennecott Copper Corporation | Hydrometallurgical brass dust reclamation |
| ES484721A1 (en) * | 1978-09-25 | 1980-09-01 | Texas Gulf Inc | Treatment of zinc plant leach residues for recovery of the contained metal values. |
| JPS58141346A (en) * | 1982-02-15 | 1983-08-22 | Dowa Mining Co Ltd | Collecting method of metal from industrial by-product |
| CA1216157A (en) * | 1982-09-29 | 1987-01-06 | Donald R. Weir | Process for separately recovering zinc and lead values from zinc and lead containing sulphidic ore |
| LU85385A1 (en) * | 1984-05-28 | 1986-01-29 | Mines Fond Zinc Vieille | PROCESS FOR LEACHING SULPHIDES CONTAINING ZINC AND IRON |
| DE3634359A1 (en) * | 1986-10-09 | 1988-04-21 | Ruhr Zink Gmbh | METHOD FOR PROCESSING RESIDUES FROM HYDROMETALLURGICAL ZINC PRODUCTION |
| US4778520A (en) * | 1987-03-26 | 1988-10-18 | University Of Waterloo | Process for leaching zinc from partially desulfurized zinc concentrates by sulfuric acid |
-
1989
- 1989-12-15 GB GB8928368A patent/GB8928368D0/en active Pending
-
1990
- 1990-12-14 AU AU69131/91A patent/AU6913191A/en not_active Abandoned
- 1990-12-14 CA CA 2076025 patent/CA2076025A1/en not_active Abandoned
- 1990-12-14 WO PCT/CA1990/000442 patent/WO1991009146A1/en active Application Filing
- 1990-12-17 ZA ZA9010122A patent/ZA9010122B/en unknown
-
1992
- 1992-08-14 US US07/929,927 patent/US5348713A/en not_active Expired - Fee Related
Also Published As
| Publication number | Publication date |
|---|---|
| AU6913191A (en) | 1991-07-18 |
| US5348713A (en) | 1994-09-20 |
| WO1991009146A1 (en) | 1991-06-27 |
| GB8928368D0 (en) | 1990-02-21 |
| ZA9010122B (en) | 1991-10-30 |
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